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Access to the Rosebery Orebody, which dips into the western slopes of Mount Black at 45¦, has been achieved in sequence by the following means: - A series of adits down to 8 level - An inclined shaft (No. 1 Shaft) sunk internally from 8 level to 14 level - A vertical circular concrete lined shaft (No. 2 Shaft) sunk internally from 4 level to 19 level. Although the orebody has been fairly well defined to 19 level, deeper ore occurrences have been outlined by a series of drill holes decreasing in frequency with depth to 26 level. No. 2 Shaft (sink) was stopped some 10 m below 19 level in order to make an early start on doubled production at West Coast Mines, with the intention of subsequently extending the shaft or of driving a decline from the lowest operating horizon for deeper access. The decline approach was adopted some six years after production commenced from No. 2 Shaft. The reasons for that decision are outlined, together with an account of the constraints and design requirements imposed by commencing by a vertical shaft in the hanging wall of the orebody.

Jan 1, 1982

By William R. Sharp

This paper describes a graphics station and graphics software designed specifically for use within the minerals industry. Through the graphics station, a user is able to recall maps for modification, display them on a color CRT, send the work to a plotter, and restore the work when finished. The engineer may use the system for the storing of data, per- forming engineering calculations, doing word processing, and performing general file maintenance. In addition, the system may be used to communicate with a mainframe computer for such purposes as: transmission of large data files, graphic 3-dimensional rotation of a mineral deposit or mine design, contouring of spacial data, and other data processing work.

Jan 1, 1982

By Richard L. Stein

INTRODUCTION Increasing a miner's chance of surviving a disaster requires advance planning for emergencies, adequate training of personnel, and the provision of proper survival equipment. It is extremely important that fires and toxic gases be detected early, and that warning systems such as stench, audible and visual alarms, and communications systems be provided to allow timely escape. However, situations do occur where timely warning is not provided and immediate escape is not possible, regardless of the precautions that are taken. Under such circumstances, a combination of new per¬sonal protective devices and old techniques can provide protection that is adequate to save a miner's life. SELF-RESCUE EQUIPMENT Self-rescuers are devices which provide a short-term supply of respirable air under the potentially lethal con¬ditions following a mine disaster. Functioning either as filtration devices for the elimination of carbon monoxide (CO) or as oxygen supplies, self-rescuers allow an endangered miner a limited amount of time in which to effect an escape or reach a place of temporary safety. Carbon-Monoxide Self-Rescuers For almost 50 years, filtration self-rescuers have protected miners' lives following explosions or fires. Designed to eliminate carbon monoxide from inhaled air, the first of these devices was approved in the 1920s and has evolved into the two units that are approved today-the Mine Safety Appliance (MSA) W-65 and the Draeger 810. Both of these units rely on the cata¬lytic conversion of very toxic carbon monoxide to rela¬tively safe levels of carbon dioxide (CO,). As shown in Figs. 1 and 2, both of these filtration self-rescuers operate in the same manner. On inhalation, the mine air passes through both coarse and fine filters that remove the dust, preventing the dust from coating the chemical beds or entering the miner's mouth. The air then passes over a drying agent that removes water vapor; this is required to prevent the water vapor from poisoning the catalyst. Subsequently, the dried air passes over the Hopcalite catalyst, where carbon monoxide is converted to carbon dioxide. Then, the air passes through a heat exchanger that reduces the temperature of the inhaled air. Finally, the air is inhaled by the user. Air exhaled by the user passes through the heat ex¬changer and exits the device through an expiratory valve. Both units of this type are subject to certain limita¬tions: 1) The units do not protect the user against oxygen¬deficient air; when air is inhaled through this type of self-rescuer, the device only removes the carbon mon¬oxide. If the mine air contains less than 15% oxygen, anoxia is inevitable. Symptoms of anoxia include dizzi¬ness, shortness of breath, quickened pulse, and deeper and more rapid respiration while the victim is at rest. During heavy exertion such as would be expected dur¬ ing escape efforts, a 15% oxygen level can cause loss of consciousness. 2) The units do not protect against excessive levels of carbon dioxide. Available data on carbon monoxide and carbon dioxide concentrations following explosions or fires are limited, but there are indications that the carbon monoxide concentration can rise to 2% and the carbon dioxide concentration can reach 5 or 6% of the mine air (by volume). In some situations, the in¬haled air could contain up to 6 or 7% carbon dioxide. At a carbon dioxide concentration higher than 2%, breathing patterns can be affected adversely. At a con¬centration of 6 or 7%, the effects include severe respira¬tory distress, with unconsciousness resulting from exer¬tion such as that needed for escape activities. 3) The units do not protect against high inhalation temperatures. The catalytic oxidation of carbon mon¬oxide to carbon dioxide is an exothermic reaction that evolves a large amount of heat. As the carbon monoxide concentration increases, the temperature of the inhaled air also increases. Available data show that above 1.5% carbon monoxide, the air inhaled through some self¬rescuers can be as high as 90°C (194°F). At those

Jan 1, 1982

By P. E. Weidmark

"PART 1-Published PapersAllcott, G.A., and Archibald, D.E.; ""Description of pillar behaviour at Heath Steele Mines""; Can Min Metall Bull; 74:834:80-87; 1981.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of rock Mechanics; 3:32-33; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:37-38; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:38; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses"" ; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:41; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:42; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:46; 1979.Baar, C.A.; Discussion of Theme 2: ""Tectonophysics""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:60; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3: 116; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:116-117; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:117; 1979.Baar, C.A. '; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:118-119; 1979.Baar, C.A.; Discussions (2) of Theme 4: ""Underground openings""; Proceedings 3rd Congress lntnl Soc of Rock Mechanics; 3:119; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:120-121; 1979."

Jan 1, 1982

By G. B. Atkinson

To develop technology to help assure an ample domestic supply of zinc, the Bureau of Mines is investigating an aqueous chlorine-oxygen leaching process for treating complex sulfide concentrates. This report discusses the purification of aqueous C12-02 leaching solutions by utilizing zinc cementation to remove Cd, Co, Cu, Ni, and Pb. Additional processing of the purified solution is necessary to produce ZnC12 crystals suitable for molten-salt zinc electrowinning cells. The effects of solution temperature, reaction time, zinc dust concentration, ZnCl2 concentration, and degree of agitation on the removal of impurities were investigated for both simulated and actual leaching solutions. Zinc dust addition to a leaching solution at 30° C decreased the Cd, Cu, Ni, and Pb concentrations to acceptable levels of less than 3 mg/1 each. By increasing the solution temperature to 75° C, the cobalt content was decreased to less than 5 mg/l, an acceptable level for producing electrolytic cell-grade ZnC12. Impurity removal rates from ZnCl2 solutions are Pb>Cu=Cd>Ni>>Co. An increase in the solution temperature increases the rate of removal for these metal impurities.

Jan 1, 1982

By Gordon Lewis

A complex mixture of hydrated reaction products is obtained from hydrothermally treated refractory cements and castable concretes. TGA can be used both to identify the product phases in the mixture and to determine a quantitative relationship among the phases. When reacted mixtures contain compounds that decompose simultaneously, or nearly so, additional assumptions must be introduced in order to develop a quantitative phase analysis. It is the purpose of this paper to examine some of these assumptions and to demonstrate their application, particularly to refractory cement systems.

Jan 1, 1982

By Jerry W. Stengel

The Bureau of Mines subjected two manufacturers' chemical oxygen (K02) self-rescue breathing apparatus to a series of laboratory environmental treatments designed to simulate various conditions in underground coal mines. The environmental treatments consisted of extremes of temperature, shock, and vibration. The tests were designed to be used as predictors of the ability of the self-rescuers to survive those environmental insults with no degradation in their protection to the wearer. Based on the severity of the treatments, simulating conditions more severe than offered by the mining environment, the two apparatus tested should be able to withstand the abuse offered by the mining environment and still function as intended in an emergency. Correlation of these tests with results of long-term field evaluations is needed to provide confidence in the laboratory tests as predictors.

Jan 1, 1982

By Wm. Terry Siemers, Frank E. Kottlowski, George S. Austin

The value of industrial rocks and minerals produced in New Mexico in 1978 exceeded $227 million. The state ranked first in the nation in production of potash and perlite and was a leader in pumice production. Other commodities with sizable production in New Mexico include barite, caliche, clay, dimension stone, fuorspar, gypsum, humate, limestone, mica, salt, sand and gravel, scoria, and silica sand. Non-metallic resources are distributed throughout the state, but development and exploitation is concentrated in western and southeastern New Mexico. The value of industrial rocks and minerals within the state grew about 9% to more than $227 million in 1978. Continued growth, however, depends on continued demand, exploration, and development.

Jan 1, 1982

By Asheim, Harald A.

Parameter estimates affect both the optimum field development decisions and the economic result. A general methodology for assessment of the effects of parameter uncertainty in offshore development decision situations is proposed, and used to analyze aspects of the initial development situation for the Statfjord field.

Jan 1, 1982

Existing fluid flow techniques may be readily applied to the design of ranges for seam gas drainage. The Reynolds number characterises the flow regime within the pipe and it can be used in conjunction with data on absolute pipe surface roughness to determine the system friction factor. The application of the solution to the isothermal compressible flow equation utili- sing Moody friction factors and the concept of equivalent length of pipe for fittings and valves provides a practical method for sizing gas ranges. The physical capacity of a range can be decreased by the existence of significant quantities of water in the range and also by the ingress of air into the system at poor pipe joints etc. Both problems can be elimin- ated by adequate attention to design detail.

Jan 1, 1982

Detailed observations were made for two existing continuous miner and shuttlecar systems. A simple model for the shift average cycle consisted of three time components - delays, mining and transport. Cycle improvement and productivity improvement factors are derived and used to predict the theoretical improvement which the observed cycles could achieve. Various strategies of action ranging from short-term optimisation through to long-term planning for continuous haulage systems are evaluated. High potential for improvement is shown for a wide range of operating conditions. An overview is presented for systems, currently available or under active development, which can accept coal directly from continuous miners and offer continuous haulage to the surface. Systems are broadly classified as mechanical-roof mounted, mechanical-floor mounted and hydraulic transportation.

Jan 1, 1982

Western Mining Corporation, managing partner of the Windarra Nickel Project are in- stalling a raise-bored haulage shaft and under- ground crushing system at the Mt. Windarra Nickel Mine. The system will be the second of its type installed by W.M.C., the first being in- stalled at the Otter-Juan mine in Kambalda in 1976/77. Both haulage systems became necessary as a result of the increasing production limi- tations, and escalating costs of a deep decline mine. The Windarra haulage system required a single 3m diameter vertical raise drill hole to a depth of 470 metres. Two five tonne skips in balance will hoist crushed ore and mullock to the surface, at a design rate of 200 tonnes per hour, effectively reducing truck haulage leads by up to 50%. It is expected the project will be in full operation by mid 1982 and will enable the mine to achieve an annual production rate of 600,000 t.p.a. by 1983.

Jan 1, 1982

By H. C. Heard

INTRODUCTION As the world's population expands and nations struggle to better their relative position and standard of living, increased emphasis is being directed to the exploration and production of mineral resources, to the mitigation of natural hazards such as earthquakes and volcanic eruptions, and to geotechnical engineering projects such as high-level nuclear waste disposal, underground military facilities, and transportation tunnels. Several topics are common to all of these endeavors. One is the characterization of the rock mass in situ, including the definition of the initial environment (e.g., temperature, stress, etc.). A second is, depending upon the use of the underground or the desired result, the manner in which the boundary conditions or the environment are to be changed by man. Next, the physicochemical behavior of the rock mass must be measured, or inferred with uncertainty limits placed on this behavior, over the appropriate range of environmental conditions (e.g., temperature, chemistry, etc.) for the application proposed. And last, an appropriate, validated predictive model must be applied to estimate the rock mass response to the perturbation. In the period preceding about 1955, only relatively simple testing of soils and rocks was commonplace, partly because the environment (depth) to which these data were applied was not so extreme as it is at present and partly because less precise predictions of material response were acceptable. Only limited and relatively simple test data seemed to be needed to guide the engineering. At the present time and certainly in the future, as we extend our interaction with the earth's crust to greater depths and impose more extreme perturbations on it, we need a larger body of data, deter- mined under much more complex conditions, in order to make the required predictions. The present societal problem of high-level radioactive waste disposal is a case in point. We in the rock

Jan 1, 1982

By S. R. Gilbert, B. W. Burdett, L. B. Hales

INTRODUCTION Grinding control is currently the most publicized control subject in the minerals industry. Flotation control and crushing control have not as yet received the widespread attention that grinding control has. The primary reasons for this center around the high capital costs of grinding equipment and the fact that grinding mills are extremely energy inefficient. In a typical copper concentrator the energy requirements to produce refined copper are: (1) [Total Energy Distribution (Kw-hr/yr)% Mining4.93 x 10817.3 Mineral Processing9.63 x 10833.8 Smelting and Refining13.9 x 10848.9] One-third of the total energy used for copper production is for mineral processing. A breakdown of this energy into process steps yields the following: [Total Energy Distribution (Kw-hr/yr)% Crushing and Grinding 7.16 x 108 73.8 Flotation 0.76 x 108 7.8 Materials Handling1.78 x 10818.4] It is apparent that comminution operations consume the largest proportion of the energy for processing and as such become major targets for improvements in efficiency. Unfortunately, size reduction operations are extremely inefficient. Efficiency calculations for grinding range from less than 0.1% to about 5% depending on the evaluation criteria. Because of these facts, grinding control has been and will continue to be the focus of major process control efforts. OBJECTIVES AND BENEFITS All benefits achieved as the result of grinding control are derived from modifying physical parameters which improve the efficiency of the grinding mill. These parameters include:

Jan 1, 1982

By James E. Kolesar

An essential prerequisite to expanding the application of longwall mining techniques in the United States is the development of a better understanding of the mechanics of rock mass deformation near the mined seam, in the inter- burden and at the surface. The results of studies of this type will provide a basis for the development of empirical and/or numerical modeling techniques to predict the rock mass response to mining both near and far from the mined seam. This will lead to improved gate and panel design, resulting in maximum recovery, increased productivity, as well as provide a mechanism for assessing likely surface effects induced by long- wall mining. Presented here is the instrument plan to be used at Kitt Mine No. 1 in the study of rock mass response to two adjacent longwall panels.

Jan 1, 1982

By H. Wagner

INTRODUCTION One of the most important tasks facing the opera¬tors of deep-level gold mines is supporting the roof strata in the immediate vicinity of the stope faces. Some of the deeper gold mines operate at depths ex¬ceeding 3.0 km; the extreme mining depths and the tabular nature of the gold-bearing reefs result in high stress concentrations at the abutments of the stope faces. These, in turn, lead to extensive fracturing of even the hardest quartzites. Although most of these fractures develop in a stable manner, there have been many instances where rock failure occurred suddenly, resulting in a considerable amount of damage to the underground workings. Such violent rock failures commonly are known as "rock bursts," and they are typical of many hard-rock mines that are operating at great depths. To safeguard mining personnel and equipment against the effects of stable and unstable rock failures, adequate support must be provided to the fractured roof as quickly and as effec¬tively as possible. Traditionally, timber props and timber packs have been utilized to provide roof support in gold mine stopes up to 1.5 m wide. Both methods of support suffer from the fundamental disadvantage of generating sup¬port only in reaction to movement of the strata. Since unrestrained strata movement often results in a break¬down of the frictional forces that preserve the stability of the intensely fractured roof strata, it is essential that stope supports have rapid load-bearing characteristics. An equally important requirement is that the roof sup¬ports be able to accommodate displacements of the rock mass. A third requirement of roof supports in gold mine stopes is that they must be able to accommodate the high rates of stope convergence typical of rock bursts. The hydraulic prop is the obvious choice for sup¬porting the stope faces in deep-level hard-rock mines. Since hydraulic props can be installed with a positive setting load, they can provide roof support that is independent of strata movement. By equipping the hydraulic props with yield valves, they meet the require¬ment of being able to accommodate rock-mass displace¬ments. However, the requirement of being able to accommodate high rates of stope convergence without overloading or damaging the rock strata was rather difficult to meet. The first introduction of hydraulic roof supports into gold mine stopes showed that conventional hy¬draulic props, such as those used in coal mines, were completely unsuitable for a number of reasons, in¬cluding: 1) Since drilling and blasting are used to break the ore in gold mine stopes, supports installed close enough to be effective are exposed to the vigor of the blast. Conventional coal-mining props suffered extensive blast damage, despite being protected by a blast barricade. 2) The environmental conditions in gold mine stopes favor corrosion. The high quartz content and the abrasiveness of the ore further aggravate the problem by damaging the protective coatings on the hydraulic surfaces. Simply changing the design of gold mine props to have the piston project from the bottom rather than from the top resulted in dramatic improvements in the service life of the props. The improvement was attributable to external water in the stope washing the fine quartz grit away from the hydraulic surfaces, rather than concentrating the particles between the cylinder and piston of the prop. 3) The most obvious deficiency of conventional coal mining props was their inability to yield rapidly enough to accommodate the high convergence rates of rock bursts without overloading the supports. As a result of the first disappointing results with hydraulic supports in narrow hard-rock stopes mined by drilling and blasting, a comprehensive research and development program was undertaken. The hydraulic props resulting from several years of research and exten¬sive field trials are described herein. To date (1978), more than 130,000 of these props have been manufac¬tured and used successfully in a number of mines and under a variety of geological conditions. Although many of the detailed specifications are directed spe¬cifically to gold mining conditions, there are a number of points that are of general importance in the design of underground support systems for tabular excavations in hard-rock mines of any type. SPECIFICATIONS FOR RAPID-YIELDING PROPS Yield Requirements The yield requirements for hydraulic props in deep¬level hard-rock gold mines are summarized in the fol¬lowing paragraphs. Slow Yield Load: The design of the yield valve must allow the yield load to be preset to between 300 and 400 kN. This flexibility is necessary to account for the differences in the strength properties of the roof and floor strata. Drop in Load Under Slow Yield: The drop in sup¬port load during a slow yield should be as small as possible, but it should not exceed 5% of the slow-yield load. Rapid-Yield Load: The load exerted by the prop during periods of rapid stope convergence should be as close as possible to the slow-yield load. The rapid-yield load should not exceed the slow-yield load by more than 15% at yield rates of 1.0 m/s. The limitation of the upper valve of the yield load is important; it deter¬mines the hydraulic performance of the yield valve. The hydraulic props that satisfy this criterion have performed well, even under severe rock bursts. Load Changes Under Rapid Load: On some of the earlier rapid-yielding props, conditions of dynamic in¬stability were observed; these conditions resulted in a complete breakdown of support resistance. It is essen¬tial that the force oscillations during rapid yielding be as small as possible, not exceeding -*5% of the average rapid-yield load at a yield rate of 1.0 m/s. Longitudinal Resilience: During rock bursts, vibra¬tions of the roof and floor sometimes are observed. To

Jan 1, 1982

By Frank L. Marley

Mexican silver mining operations have existed prior to the sixteenth century. Recovery processes throughout the centuries were developed at the many mining localities and have progressed through smelting, the patio amalgamation process to cyanidation and flotation. A study of the present silver mining zones is reported in which the ores are evaluated in terms of the mineralogical constituents, their state in the matrix and the effects these are noted to impose upon the efficiency of recovery. Silver mineral identifications were by quantitative ore microscopic and scanning electron microscopic energy dispersive X-ray analysis.

Jan 1, 1982

By D. Christofferson

As an alternative power source to eliminate trailing cables, ESD Corporation is installing a flywheel system on a shuttle car. Flywheel power is extracted by a high-speed generator, developed by Lear Siegler Corp,, and the system -- including vacuum and lubrication and cooling auxiliaries -- is supplied by the aerospace group, Rocketdyne Division of Rockwell International. The flywheel system is being developed under a three-year contract with the Department of Energy. It is a follow-up contract to a previous DOE-contracted study by General Electric Co. in which it was shown that a flywheel can be a feasible power source for a shuttle car from technical, economic, and safety standpoints. This paper addresses the status of the program, design decisions established, and plans for laboratory test procedures and underground demonstration. The flywheel-powered shuttle car is a conventional 6-ton capacity shuttle car equipped with a 7-rotor flywheel energy storage package. The energy package is the sole source of power for the shuttle car through a complete mission duty cycle (MDC) of tramming to the miner, receiving a load of coal, and returning to the secondary haulage system. Energy stored in the system exceeds MDC requirements and waiting times to provide an ample margin of reserve energy. The energy storage system is recharged by spinning the rotors back up to maximum speed each time the shuttle car comes back to the secondary haulage system. The electrical charge motor is connected to a power source to recharge the system while the load of coal is being discharged.

Jan 1, 1982

By C. C. Hatch, C. R. Larsen, A. L. Mular

INTRODUCTION As part of a general project to simulate a concentrator, the Brenda Mines Limited secondary crushing plant, located near Peachland, B. C., was sampled during the summer/winter of 1975 and during the summer of 1978 to develop a digital crushing plant simulator specific to the Brenda operation. The purpose of this simulator was to study the influence of manipulable process variables on important responses, to permit study of alternative circuit configurations if desired and to perform off-line optimization. This chapter describes the methodology employed to develop the simulator (1,2) along with typical applications thereof. SECONDARY CRUSHING PLANT FLOWSHEET Figure 1 is a schematic diagram of the Brenda Mines Ltd. secondary crushing plant circuit. The 41,000 tonne coarse ore stockpile(a) is -17.8 cm oversize (3B belt) from two 2.4 m x 6.1 m double deck vibrating screens (not shown). The screens are fed by 2 conveyor belts operated in series with two 198 cm apron feeders that handle discharge from a 152 cm x 226 cm primary gyratory crusher(520 KW). The primary crusher (not shown) treats up to 2725 tonne/hr of run of mine ore. Minus 1.9 cm undersize from the primary screens (#19 belt), referred to as the primary fines, joins secondary screen undersize (#18 belt) for conveyance to fine ore bins (not shown). Feed from the coarse ore stockpile is transferred by means of vibrating feeders and two 120 cm conveyors (#6 and #7) to transfer belts #8 and #9 (Figure 1). The belts feed two secondary 2.1 m standard heavy duty cone crushers (bl and b2). Each crusher is equipped with a modified coarse mantle liner and a standard coarse bowl liner. Mantle throw is 9.2 cm; nominal close side set is 3.2 cm. Crusher .countershafts (443 rpm) are driven by single 260 KW synchronous motors (710 rpm). Discharge from the secondary cone crushers joins discharge from four 2.1 m tertiary heavy duty shorthead cone crushers (fl, f2, f3 and f4). With the exception of the cavity (modified medium mantle liner and standard coarse bowl liner; close side set is nominal .84 cm) and motor (224 KW), these machines are similar to the secondary crushers. The combined discharge is transferred from a 183 cm conveyor to #14 belt which empties into a 1452 tonne secondary screen surge bin (c). Vibrating feeders beneath the bins feed five 2.4 m x 6.1 m double deck vibrating screens (dl, d2, d3, d4 and d5) sloped at 200 from the horizontal. The upper decks are punched plates with 2.7 cm x 7.6 cm slots. Lower decks are woven wire (1.27 cm diameter) sections with

Jan 1, 1982

By C. H. Cheh

The amount of fines in run-of-mine coal has been increasing and coal quality has been deteriorating as a result of the increased use of mechanized and continuous mining methods. Removing the mineral matter and pyritic sulfur from coal prior to combustion will lower transportation costs, improve power plant performance and reduce sulfur dioxide emissions. Since physical cleaning methods can only remove individual particles of mineral matter and pyrite, it is necessary to crush and grind coal to fine sizes to separate these contaminants so they can be removed by deep cleaning processes. The increased portion of fines in run-of-mine coal together with attempts to increase the mineral matter removal efficiency have stimulated interest in fine coal cleaning technology. However, there are many problems to be solved before fine coal cleaning can be applied to power plant coals. These problems include: 1. coal cleaning plant capacity is reduced when conventional cleaning circuits are modifies to deep clean coals; 2. existing equipment and technology is not suitable for cleaning fine coal; 3. coal recovery is reduced; and 4. fine coal is difficult to dewater, dry, handle, transport and store.

Jan 1, 1982